v118n11a10 structural characteristics of strata overlying ... · longwall face (sgmhlw). ><...

The coal seams in the Shendong coalfield,located in the northwest of China, arecharacterized by low dips and largethicknesses and are thus suited for fullthickness extraction. Since the mid-1990s, theShendong coalfield has been developed intoone of the largest in the world (Ning, Liu, andTan. 2014; Ning et al., 2017). Recentdevelopments in mining equipment have madeit possible extract entire seams up to 5.0 mthick in a single cut. This method, termed the‘super-great mining height longwall face’(SGMHLW) method, has been widely adoptedin the Shendong coalfield (Table I) (Ju and Xu,2013; Peng, Li, and Zhou 2015; Zhang, Fan,and Ma 2011). According to some studies,sudden support-closure incidents often occurwhen using the SGMHLW method. The key tosolving this engineering issue is to have aclear understanding of the structuralcharacteristics and movement laws of theoverlying strata. For instance, the load on thesupports is related to the movement of the roofstratum as a response to the mining activities.

To date, extensive studies have beenconducted to obtain a clear understanding of

the mining-induced cave-in response of thestrata (e.g. roof movement and failureresponses). These studies have used differentmethods, such as numerical simulations,theoretical analyses, physical modelling, andon-site investigations (Ghose and Dutta, 1987;Yasitli and Unver 2005; Trueman, Lyman, andCocker 2009; Shabanimashcool, Jing, and Li2014; Tan, Li, and Ning 2017; Jiang, Sainoki,and Mitri 2017). Based on field observations,Peng and Chiang (1984) suggested that thefirst cave-in event involves shear fracture ofthe main roof before that of the face, while thesubsequent and periodic cave-in eventsinvolve cantilever instead of Voussoir beamcollapse. Song (1988) proposed that the rockblocks in the main roof stratum rotate andinterlock with each other to form a jointedVoussoir beam when the stratum deflectsdownwards. Sofianos (1996) developed amodel of Voussoir beams in the hard roofthrough numerical simulations. These studiesindicated that roof cave-in is a dynamicprocess involving rock fracturing,disintegration, and movement. However, allthese studies were conducted with a miningheight of less than 6.0 m, i.e. under normalmining height extraction conditions.

In recent years, considerable attention hasbeen given to the strata failure and movementpattern induced by the mining of thick coalseams, for example, the longwall top coalcaving (LTCC) and SGMHLW methods (Gongand Jin 2008; Yu, Zhao, and Kuang 2015; Yu,Zhao, and Xiao 2017). Owing to the highoutput and high efficiency of the SGMHLWmethod compared with the LTCC method atpresent, it is popular for mining thick coal

Structural characteristics of strataoverlying of a fully mechanized longwallface: a case studyby J. Wang, J. Ning, L. Jiang, J-Q. Jiang, and T. Bu

In coal mining in China, the cutting height of the shearer in longwall facesis increasing. Owing to the increase of extraction height, the caved roofstrata area is enlarged, and new issues are being encountered, such asdetermination of a suitable working resistance for the shield and control ofthe roof. Through field observation and theoretical analysis of the firstlongwall face with a height of 6.0 m in the Bayangaole mine, a three-stagestructural model was developed. Stage I is defined as the period in whichthe lower immediate roof (LIR) caves into the goaf and is broken intoirregular shapes of various sizes. Stage II is defined as the period in whichthe upper immediate roof (UIR) breaks and impacts the LIR. In stage III,the main roof breaks into blocks and then impacts the UIR. With respect tothese three stages, a suitable method was identified for calculating theworking resistance of the shield support for a super-great mining heightlongwall face (SGMHLW).

Strata behaviour, strata movement, working resistance, super-greatmining height longwall face.

* State Key Laboratory of Mining DisasterPrevention and Control Co-founded by ShandongProvince and the Ministry of Science andTechnology, Shandong University of Science andTechnology, China.

© The Southern African Institute of Mining andMetallurgy, 2018. ISSN 2225-6253. Paper receivedJul. 2017; revised paper received May 2018.

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http://dx.doi.org/10.17159/2411-9717/2018/v118n11a10

Structural characteristics of strata overlying a fully mechanized longwall face: a case study

seams in China. Recently, a significant number of studieshave been performed under SGMHLW extraction conditions.Literature reviews show that with the increase in extractionheight, the mining-induced strata failure area is enlarged andthe strata behaviour differs from that with a normalextraction height (Unver and Yasitli, 2009; Singh and Singh,1999; Alehossein and Poulsen, 2010). Owing to complexgeological conditions, roof fracturing, induced movements,and roof management are still unclear, in particular when themining height is greater than 6.0 m.

Determination of the shield capacity has been a researchobjective in many countries, resulting in a number ofmethods for its calculation (Coulthard, 1999; Verma and Deb,2013; Lawson et al., 2017; Batchle,r 2017; Prusek, Płonka,and Walentek, 2017). As early as the 1980s, mostresearchers suggested that external loading on supports wasrelated to the weight of the roof strata (Henderson, 1980).Currently, the determination of external loading on the shieldsupports is still a critical issue for roof management. Thereare four main methods for determining shield capacity: thedetached roof block method, shield-leg pressuremeasurement method, design of powered support selectionmodel, and yielding foundation model (Smart and Redfern,1986; Gilbride, Richardson, and Agapito 1998; Trueman,Lyman, and Cocker, 2009; Islavath, Deb, and Kumar, 2016).All these methods have been commonly used by supportmanufacturers and coal operators. In summary, thesetheories and methods are used for calculating loads on shieldsupports based on the strata movement and failure responsesinduced by mining in a longwall face with a mining heightless than 6.0 m. The determination of shield support capacityis still challenging because of the limited understanding ofthe behaviour of the roof strata and of the mechanismsresponsible for external support loading when the miningheight is greater than 6.0 m.

Owing to the limited studies, some issues, such as themovement of the overlying rock strata induced through theSGMHLW mining method and the unique formulae fordetermining external loadings on the supports, require morein-depth investigation and understanding. The presentanalysis is based on the mining conditions of panel 311101in the Bayangaole Colliery, China. In this study, first, thetime-weighted average resistance (TWAR) of the shieldsupport was measured to understand the roof behaviour.Secondly, field observation was used to simulate the mining-

induced overburden failures with the advance of the longwallface. Finally, a simplified theoretical model was established todescribe the structural characteristics and movement of theroof. From these studies, a mathematical formula wasdeveloped to estimate the shield working resistance.

The Bayangaole Colliery is located in the Shendong coalfield,Ordos City, Inner Mongolia Autonomous Region, China. Thecoal-bearing geological sequence has an average thickness of730 m and contains eight coal seams, five of which can bemined with a total thickness of 24 m. The mine is currentlyextracting seam No. 3 and all the panels in this mine areusing the retreat longwall extraction method.

Panel 311101, at an average depth of 614 m, was used asthe target panel for this case study. As the first longwall facein the mining area of coal seam no. 3, Panel 311101 adoptsthe SGMHLW method. Its designed cutting height is 6.0 m.The roof of the longwall face is supported by ZZ15000/16/25standing-shield hydraulic supports with a rated workingresistance of 15 000 kN (Figure 1). A total of 150 shieldsupports were used in this panel, and were numbered from 1to 150. Seam no. 3 has a mean thickness of 6.0 m and amean dip angle of 1.5°, ranging from 0° to 3°. The panel isapproximately 260 m along the dip, and 2 550 m along thestrike. The simplified stratigraphic column of this panel,obtained by core logging, is shown in Figure 2. The roof

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Table I

Bayangaole mine 311101 6.0 Medium hard 12 782 This studyBulianta mine 22303 7.0 Soft and weak 17 612 Ju et al. (2013)Bulianta mine 32301 6.1 Soft and weak 10 022 Peng et al. (2015)Halagou mine 22401-2 5.6 Soft and weak 7 587 Peng et al. (2015)Shangwan mine 51201 5.3 Medium hard 11 856 Zhang et al. (2011)Shangwan mine 12206 6.8 Soft and weak 14 652 Peng et al. (2015)Nalinmiaoer mine 62105 6.2 Soft and weak 11 207 Peng et al. (2015)Daliuta mine 22614 5 Soft and weak 9 176 Peng et al. (2015)

Note: Mean roof load was the shield support load at the onset of roof weighting

strata are composed mainly of sandy mudstone, finesandstone, and siltstone; these have a mean uniaxialcompressive strength (UCS) of over 30 MPa. The floor ismainly sandy mudstone, with a mean UCS of 45 MPa. Theoverburden rocks are designated as medium-strong type.

To measure the shield leg pressures from shield to shield andcycle to cycle, a real-time shield-leg pressure monitoringsystem (SLPMS) was utilized, as shown in Figure 3. This

SLPMS was produced by UROICAC Ltd., China. It consists ofa host, control computer, data line, power supply equipment,and substation, among other components. The analysissoftware was installed on the control computer software toautomatically calculate the TWAR of the shield supports andrecord the shield-leg pressures at one-minute intervals.

To simplify the analysis, this paper illustrates one case(shield leg pressures of the shield support no. 100) todescribe the change rule of shield leg pressures. Figure 4shows the recorded shield leg pressures of the shield supportno. 100 when the longwall face advanced 20 and 60 m. Asshown in Figure 4, when the longwall face advanced 20 m(the nonweighting period), the shield leg load wasapproximately 7 500 kN. However, when the longwall faceadvanced 60 m (weighting period begins), the shield leg loadreached the early warning value of 12 000 kN. It can bededuced that the load on the shield support increasedconsiderably because of the beginning of the weightingperiod. In the figure, the early warning value was setaccording to engineering experience (in Table I). During theperiod marked ‘overhaul’ the longwall face did not advancebecause of faulty equipment.

[1]

where Pt is the average support resistance during period ti,and Tt is the time that a mining (supporting) cycle takes.

Figure 5 shows a typical support leg resistance variationin a mining cycle. In an earlier study by Peng (2015),Equation [1] was used to calculate the TWAR. Figure 6shows the TWAR for a face advance distance of 500 m. As


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Main station

Power Substation 1#Substation 2#

Substation 100#Substation 149#Substation 150#

Substation 1# Substation 149#

Hydraulic support

Communicationsubstation

Power SupplyEquipment

Underground Control computer

Report form

Ground surfaceData cable

(a) Field layout of SLPMS

(b) Full profile of SLPMS

Host

Hydraulic support Hydraulic support

Substation 150#

the longwall face advances, three stages are observed,distinguished by changes in the slope of the curves. Thesestages correspond to stages in the movement of the roofstrata. In the following text, these three stages are named A,B, and C. At stage A, along curves ab, a'b', and a''b'', theTWAR increases slowly, and its maximum value is 3 000–3500 kN. At stage B, along the curves bc, b'c', and b''c'', theTWAR increases rapidly, and its maximum value is 4 000–5

000 kN. Interestingly, at stage C, along the curves cd, c'd',and c''d'' the TWAR shows the greatest increase: theincrement is 7 000–9 000 kN and is at least twice as large asthat in the other stages. The in situ measurementsdemonstrated that the dynamic load coefficient (DLF: theratio of shield-leg pressure during the nonweighting periodto the shield leg pressure during the weighting period (Ju andXu, 2013) presented a periodic alternation betweenshort/gentle and long/strong, and stage C was accompaniedby a large DLF, up to 2.5, which indicated a violentmovement of the overburden.

A deep hole multiposition extensometer (DMPX), producedby UROICAC Ltd., China, was used to identify the longwallmining-induced strata movement. The DMPX is a speciallyconstructed instrument that measures the differentmovements of selected overburdened strata layers in aborehole relative to a fixed point. In addition, a digitalpanoramic imaging device (DPID; UROICAC Ltd, China) wasused to determine the longwall mining-induced strata failure.The DPID is a specially constructed instrument used forinspecting roof strata conditions, such as mining-inducedmacro-fractures, inside a borehole. These monitoring deviceswere also used by Ning et al. (2017) to investigate mining-induced overburden failures.

The DMPX was applied at a borehole, and four anchorswere installed to monitor the overburden movement. Figure 7shows the locations of the anchors. The displacementmeasured at the borehole with the advance of panel 311101is shown in Figure 8. Hi (i = 1, 2, 3, 4) is defined as thedisplacement of the rock strata at anchor i. The horizontaldistances from panel 20107 to the boreholes is denoted as D,which is taken as negative when the working face does notadvance across the location of boreholes and positive when


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the working face advances across the location of boreholes.As shown in Figure 8, when D = 0 m < D < 20 m, theoverlying strata were observed to move. When D = 38.2 m,H1 rapidly increased to 986 mm, which was approximatelyone-sixth of the thickness of coal seam no. 3. This suggeststhat the strata monitored by anchor no. 1 moved as one unit.When D = 40 < D < 50 m, H2, H3, and H4 began to similarlyincrease. When D was approximately 60 m, H2, H3, and H4increased rapidly and reached a maximum of approximately696, 713, and 709 mm, respectively. This suggests that thestrata monitored by anchors no. 2, 3, and 4 moved as oneunit when the longwall face was approximately 40–60 mfrom the borehole.

The above analysis was validated through the monitoringof mining-induced macro-fractures (the DPID was used tomeasure the magnitudes of fractures, voids, and dislocationsin the borehole wall). Figure 9a shows the fractures in theborehole wall recorded 7–8 m above the top of the coal seamfor D = 40 m. It could be deduced that when panel 311101advanced 38.5 m, the strata monitored by anchor no. 1 (fine

sandstone) fractured. At this time, the working surfacehydraulic support pressure increased, leading to roofweighting. This implies that the fracturing of fine sandstonesled to roof weighting. Similarly, when D = 60 m, fractureswere observed in the borehole wall 15–41 m above the top ofthe coal seam (Figures 9b–d), indicating that the stratamonitored by anchors no. 2, 3, and 4 were fractured as oneunit. At this time, roof weighting also occurred.

In brief, the field monitoring provided valuable and reliabledata, indicating that the strata behaviour induced bySGMHLW is very violent and the dynamic load is much largerthan when mining thinner seams. Owing to the considerableextraction height, the mining-induced strata failure area isenlarged and the progressive failure can be divided intodifferent stages. This is described in the following sections.

To understand the progressive failure of the overlying roofinduced by the SGMHLW operation, a simplified theoreticalmodel was developed based on the field measurements andcantilever beams (Diederichs and Kaiser, 1999). Hereafter,model A represents the first occurrence of roof weighting andmodel B the subsequent periodic roof weighting.

Based on the field measurements and theoretical research,the progressive failure of overlying strata in model A isdivided into three stages:

� Stage I: In this stage, the lower immediate roof (LIR)caves into the goaf and is broken into irregular shapesof various sizes. LIR refers to the weak overlying layerwhich is located above the coal seam and has a similarthickness as the seam. As the longwall face advancesfrom the set-up entry, the LIR bends and sags. Whenthe face moves beyond the critical span of the LIR, it


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breaks into fragments of different sizes and then cavesinto the goaf, as shown in Figure 10.

� Stage II: In this stage, the upper immediate roof (UIR)is broken and impacts the LIR. As the face continues toadvance, the UIR sags downward. Once the longwallface moves beyond its critical span, the UIR is fracturedand split into two rock beams (A and B), as shown inFigure 11. These two rock beams rotate with respect toeach other to form a Voussoir beam above the mined-out void. Owing to the large mining height, the panelvoid cannot be completely filled by the caved-inmaterials. The Voussoir beam continues to deflectdownwards, ultimately resulting in a buckling failure.At that moment, rock beam B impacts the shieldsupport via the LIR while rock beam A falls directly intothe caved zone. The cave-in of the UIR involves similarprocesses to those in the conventional immediate roofthat directly caved into the gob; however, it alsoinvolves instability before the balance of the Voussoirbeam is broken. Owing to this special property, thedynamic impacts of the UIR cave-in event are harmfulto the stability of the longwall shield supports.

� Stage III: In this stage, the main roof is broken intoblocks and then impacts the UIR. Once theadvancement of the working face has reached thecritical span of the main roof (Figure 12), the firstcave-in occurs as the main roof is broken into two rockbeams owing to the occurrence of tensile fractures at itsmid-span. When the main roof fractures, the brokenroof beam quickly sags, generating dynamic impactsbecause of the large bed separation between the mainroof and UIR. At this stage, owing to the impact forceinduced by the sudden sagging of the main roof, thelongwall support load immediately increases, increasingthe dynamic load coefficient. The caved height in theroof is the highest in the middle of the caved zone,where the panel void is fully filled by the caved-inmaterials.

Field experience shows that after the first roof weighting, thelongwall face enters into the second phase of overburdenmovement. In this phase, the main roof breaks periodically

behind the face after every critical advance so that a periodicweighting is placed on the longwall shields. The periodiccave-in is a progressive process which starts when theexcavation reaches a critical length. During each occurrenceof model B, the overburden movement can also be dividedinto three stages. During stage I, the LIR caves immediatelybehind the longwall supports. During stage II, the UIR hangsin the form of cantilever beams and is periodically broken atcertain intervals. In stage III, shear fracture occurs in themain roof ahead of the face, and dynamic impacts aregenerated (Figure 13).

At the beginning of stage I, both the LIR and main roofdeform gently, and thus maintain stability. Then, the LIRcaves into the goaf following the advance of the shieldsupports. Therefore, the LIR cannot maintain its self-supporting stability. In this case, at least half of the LIRweight needs to be supported by the shield canopy (Ning etal., 2014). Consequently, the equation for calculatingworking resistance for stage I can be expressed as follows:

[2]

where mz and Lz are the thickness and critical span of the LIR(m), respectively, z is the unit weight of the LIR (kN.m-3), Lkis the shield canopy length (m), and P1 is the calculatedworking resistance of the shield support for stage I.

At this stage, the main roof continues to deform gently andits movement has little influence on the UIR movement.Consequently, the force acting between the main roof and LIRcan be ignored. Therefore, the roof load applied to the shieldsupport is composed of two parts: the acting force on thesupport induced by the LIR, Q1, and the impact load inducedby the movement of the UIR, Q2. As suggested by Song(1988), Q1 can be expressed as follows:

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[3]

where Ls is the rear overhang of the LIR (m).To identify the impact force induced by the movement of

the UIR, Q2, the yield foundation model was introduced. Inthis simplified model, the shield support is regarded as aspring (as shown in Figure 14). In Figure 14a, o1 is regardedas the centre of gravity of beam B. GB and GA are regarded asthe weight of beams B and A, respectively. Here, status Arepresents the beginning of the sagging of the UIR. Status Brepresents the occurrence of instability in the Voussoir beam.Moreover, once status B occurs, the LIR is suddenly impactedand the impact load is transferred to the longwall supportsvia the Voussoir beam. Status C represents the end of impactin the LIR. Accordingly, in this study, we assumed that (i) atthe time of statuses B and C, the velocity of the UIR tended tobe zero; (ii) the UIR is regarded as a rigid body with verticalmotion, and it transmits impact force to the shield support;(iii) the shield is regarded as a linear spring; and (iv) there isno energy dissipation during the UIR movement.

At statuses B and C, the UIR velocity is zero. Thus, thekinetic energy variation is also zero. The potential energychange of rock beam B Esp, UIR Exp, and shield supportEv, can be expressed as follows:

[4]

where GB is the weight of the rock beam B (kN), h0 is theaverage displacement of the LIR covered by the shield canopy(m), x1 is the horizontal distance from the UIR beam end-fracturing to the longwall face, Lz is the first critical span ofthe UIR (m), and Fsd is the impact force induced by the UIRmovement (kN).

At this stage, the potential energy of gravity istransformed into the elastic potential energy of the shieldsupport. According to the energy conservation law (David,Robert, and Jearl, 2014), the relationship is written asfollows:

[5]

Based on Equations [4] and [5], the impact force inducedby the UIR movement can be established as follows:

[6]

The impact force transmitted to the shield support by theLIR can be expressed as follows:

[7]

where mz' is the UIR thickness (m), z' is the unit weightof the UIR (kN.m-3), and is the load transmitted coefficient,which can be obtained from Ning, Liu, and Tan (2014).

At this stage, when the LIR rotation reaches themaximum allowable sagging, the movement of the UIRinduces the impact force, namely:

[8]

where KA is the bulk factor of the LIR.At this stage, the shield support external load can be

expressed as follows:

[9]

At this stage, the working resistance P3 should be calculatedfrom three factors: The first factor is the weight of the LIR,that is, Q1; the second is load force induced by the UIRmovement, that is, Q3. The third is the impact force inducedby the movement of the main roof, that is, Q4. In this case,the weight of the LIR, Q1, can be calculated using Equation[3]. The load force Q3 induced by the movement of the UIRshould be determined from two factors: one being the weightof the UIR over the shield canopy area, and the other theweight of the UIR overhangs. As suggested by Song (1988),Q3 can be expressed as follows:

[10]

where Ls' is the rear overhang of the UIR (m).To obtain the impact force Q4 induced by the movement

of the UIR, a simplified model was also developed (Figure15). In this model, the broken main beam impacts the lowerstrata (the UIR and LIR are regarded as rigid bodies withvertical motion) rotating around point ‘o’. In Figure 15, o2 isregarded as the centre of gravity of the broken beam. InFigure 15, state D represents the downward-deflected mainroof and state E represents the occurrence of impact. State Frepresents the end of the main roof movement. At states Dand E, the velocities of the LIR, UIR, and main roof are zero.Therefore, at those states, the corresponding kinetic energy iszero.

After the main roof is broken at mid-span, thedisplacement of the roof over the shield canopy areaincreases by h1. Here, the potential energy change of themain roof ( Ejp), UIR ( Esp), LIR( Exp), and shield support ( Ev) can be expressed as follows:

[11]


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where G1 is the weight of the main roof (kN), h1 is the bedseparation between the main roof and UIR (m), L0 is the firstcritical span of the main roof (m), and h1 is thedisplacement of the roof over the shield canopy area (m).

According to the energy conservation law (David, Robert,and Jearl 2014), the impact force induced by the suddenmovement of the main roof can be obtained by

[12]

where Fjd is the impact load induced by the suddenmovement of the main roof (kN).

The impact load transmitted to the shield support by theLIR can be expressed as follows:

[13]

where mE is the thickness of the main roof (m) and E is theunit weight of the UIR (kN.m−3).

At this stage, the shield support external load can beexpressed as follows:

[14]

In model B, the overburden movement is also divided intothree stages.

At stage I, the shield support external load P4 is composed oftwo parts: the weight of the LIR over the shield canopy areaand the weight of the LIR overhangs. It can be calculatedusing Equation [3].

At this stage, the shield support external load is composed oftwo parts: the weight of the LIR (Q1) and the impact load(Q5), induced by the periodic break of the UIR. To obtain theimpact force (Q5) induced by the periodic movement of theUIR, a simplified model was developed (shown in Figure 16).In Figure 16, o3 is the weight centre. Before the UIR isbroken, the bed separation between the UIR and LIR is h2. Atthe end of the UIR movement, the displacement of the roof

over the shield canopy area increases by h0. At this stage,the potential energy change of the UIR ( Ejp), LIR ( Esp), andshield support ( Ev) can be expressed as follows:

[15]

where G2 is the LIR weight in model B (kN), Lz is the periodicbroken length of the LIR (m), and Fsd is the impact forceinduced by movement of the UIR (kN).

After the movement of the UIR, the change in potentialenergies of the UIR and LIR is transformed into elasticpotential energy of the shield support. According to theenergy conservation law (David, Robert, and Jearl. 2014), theimpact force Fsd induced by the sudden movement of theUIR can be obtained by:

[16]

The impact force Q5 transmitted to the shield support canbe expressed as:

[17]

At this stage, the shield support external load P5 can beexpressed as:

[18]

At this stage, the shield support external load P6 is composedof three parts: the weight of the LIR, Q1; the load forceinduced by the movement of the UIR, Q3; and the impactforce induced by the movement of the main roof, Q6. Theweight of the LIR, Q1, and the load force induced by themovement of the UIR, Q3, can be calculated using Equations[3] and [4], respectively.

To obtain the impact force (Q6) induced by the periodicmovement of the UIR, a simplified model was developed(Figure 17). In Figure 17, o4 is regarded as the broken mainroof. The change in potential energy of the main roof ( Ejp),UIR ( Esp), LIR ( xp), and shield support ( v) can beexpressed as:

[19]

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where G3 is the weight of the LIR in model B (kN), h1 is thebed separation between the UIR and main roof in model B(m), L1 is the periodic broken length of the main roof (m),h1 is the displacement of main roof in model B (m), x2 is

the horizontal distance from the main roof beam end-fracturing to the longwall face in model B (m), and F jd is theimpact force induced by the movement of the main roof inmodel B.

According to the energy conservation law (David, Robert,and Jearl. 2014), the impact force F jd induced by themovement of the main roof in model B can be expressed as:

[20]

The impact force Q5 transmitted to the shield support canbe expressed as:

[21]

At this stage, the shield support external load P6 can beexpressed as:

[22]

In conclusion, the abovementioned two models should beconsidered when calculating the shield support external loadwhen using the SGMHLW method. Therefore, the workingresistance of the shield support can be expressed as:

[23]

Based on the stratigraphy of panel 311101, the workingresistance of the shield supports can be calculated. As shownin Figure 1, the LIR of panel 311101 is sandy mudstone witha height of 4.0 m, which will cave and fall into the goaffollowing the advance of the coal extraction and shieldsupport. The UIR of this panel is siltstone less than 7.0 mthick. The main roof is composed mainly of fine sandstoneand sandy mudstone, and its thickness is 30.5 m. Based onexperience in mining engineering, the LIR unit weight is 22.5 kN/m3 and its bulk factor KA is 1.2. The UIR and mainroof unit weights are 22.8 and 23.8 kN/m3, respectively.

Field experience shows that cave-in events occur so thata roof weighting is placed on the longwall shields. Thevariations in response of the longwall shields can be used topredict the critical spans or breaking length of overlyingstrata. According to the analysis of the shield support TWAR(as shown in Figure 5), the increasing rate of resistance canbe divided into three main types which could refer to thebreaking of the LIR, UIR, and main roof. Therefore, thecritical spans of the LIR (Lz), UIR (Lz), and main roof (L0) inmodel A are 23.5, 38.2, and 60.0 m, respectively. The criticalspans of the UIR (Lz) and main roof (L1) in model B are 16.2and 34.7 m, respectively. A laser measuring instrument wasinstalled at the end of the support of the working face tomeasure the rear overhang of the LIR (Ls), which wasdetermined to be approximately 2.0 m. According to thesupport parameters, the shield canopy length isapproximately 5.0–6.4 m. Field investigations show that inmodel A, h1 and h1 are 0.35 and 0.20 m, respectively. Inmodel B, h0 and h1 are 0.25 and 0.65 m, respectively. Amultiposition borehole extensometer was installed in panel311101. For each extensometer, three anchors were installedin the LIR, UIR, and main roof to measure the displacements.The field measurements showed that h1 = 0.2 m and h1 =0.1 m.

According to the three types of structures that exist inpanel 311101 and the equations presented in the previoussections, the working resistance for each of the three stageswas calculated, and the results are reported in Table II. Atstage II of models A and B, the working resistances of theshield support are 4 720 and 4 517 kN, and thecorresponding dynamic load coefficients are 2.2 and 2.4,respectively. At stage III of models A and B, i.e. when themain roof undergoes violent movement, the workingresistances of the shield support are 12 782 and 10 398 kN,and the corresponding dynamic load coefficients are 2.7 and2.3, respectively. Therefore, the working resistance of theshield support with the advance of the longwall face is 1.141 MPa (12 782 kN), according to Equation [23].

Table III lists the estimated working resistance of theshield support using the equations provided by Song (1988)and Qian et al. (2010). As shown in Table III, the measuredworking resistance of panel 311101 is approximately 12 084 kN, while the working resistances estimated by usingthe equations provided by Song (1988) and Qian et al.(2010) were 7 294.1 and 8 500.4 kN, respectively, which aremuch lower than those obtained from the measurements inthis study. However, the working resistance estimated usingEquation [23] is 12 782 kN, which is very close to themeasured value. Obviously, the methods that Song (1988)and Qian et al. (2010) provide cannot be used for a SGMHLWoperation. Thus, the 6.0 m height chock shields with workingresistance of 13 000 kN were used in the next panel, that is,panel 311102 of the Gaojialiang coal mine.

In China, the SGMHLW mining method has been successfullypracticed in the Ordos coalfield. To understand the stratabehaviour induced using the SGMHLW method, an in situinvestigation and theoretical research were conducted onpanel 311101 of the Gaojialiang mine. The results presentedin this paper are summarized as follows.

1. With the advance of the longwall face, the SLPMS of


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the shield support progressed in three stages: slow,rapid, and high-speed growth. In addition, the shieldsuffered from two impact loadings induced by theoverburden movement. The numerical simulation alsoconfirmed that the progressive failure of the stratacould be divided in three stages.

2. Based on the data obtained from the in situinvestigation and numerical simulation, a simplifiedtheoretical model was built to describe the structuralcharacteristics of the overlying strata. In this model,the progressive failure of the overlying strata wasdivided into three stages. In addition, the workingresistance of the shield was determined at every stage,and then a reasonable working resistance wasidentified for the shield support under SGMHLWconditions. The reasonable working resistance isdefined as P = max (P1, P2, P3, P4, P5, P6,). Theappropriate working resistance for the 6.0 m heightchock shields of the longwall face of panel 311101 in the Bayangaole coal mine was determined at 12 782 kN.

This study was supported by National Key R&D Program ofChina 2018YFC0604703), the National Natural ScienceFoundation of China (No. 51574154, 51274133, 51474137,51574155), the Shandong Province Science and TechnologyDevelopment Plan, item (2014GSF120002), and the Tai’shanScholar Engineering Construction Fund of the ShandongProvince of China.

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1204

Table II

Stage I 2 153 1 875Stage II 4 720 4 517Stage III 12 782 10 398

Table III

Working resistance (kN) 12 084 12 782 7294.1 8 500.4Difference (%) - 5.8 –39.6 –29.7

Note: Difference with regard to working resistance measured fromSLPMS

v118n11a10 structural characteristics of strata overlying ... · longwall face (sgmhlw). ><...

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